Recovery of precious metals

ABSTRACT

A process for leaching precious metals from material containing precious metals, such as oxidic and sulfidic gold-bearing ores, is disclosed. The process includes the steps of: (i) leaching the precious metal with a leach solution containing a thiosulfate-based lixiviant; (ii) treating the material by oxidising precious metal in the material into a form that is leachable in a subsequent leaching step; and thereafter as a separate step.

[0001] The present invention relates to thiosulfate leaching of materialcontaining precious metals.

[0002] The present invention relates particularly to thiosulfateleaching of gold from gold-bearing material, such as ores andconcentrates of ores.

[0003] It is known to extract gold from ores using thiosulfate-basedlixivient systems. U.S. Pat. Nos. 4,369,061 and 4,269,622 to Kerleydescribe processes which include lixiviating with an ammoniumthiosulfate leach solution containing copper to recover gold from ores,particularly difficult-to-treat ores containing copper, arsenic,antimony, selenium, tellurium and/or manganese. U.S. Pat. No. 4,654,078to Perez et al discloses a modification of the process disclosed in U.S.Pat. No. 4,269,622 and is based on lixiviating ores with copper-ammoniumthiosulfate in a solution that is maintained at a minimum pH of 9.5.Other known processes that are based on the use of thiosulfatelixiviants include U.S. Pat. No. 5,785,736 to Thomas et al (assigned toBarrick Gold Corporation) and U.S. Pat. No. 5,354,359 to Wan et al(assigned to Newmont Gold Co).

[0004] An object of the present invention is to provide an alternativeprocess for leaching precious metals, such as gold, usingthiosulfate-based lixiviants.

[0005] According to the present invention there is provided a processfor leaching precious metals from material containing precious metals,which process includes the steps of:

[0006] (i) treating the material by oxidising precious metal in thematerial into a form that is leachable in a subsequent leaching step;and thereafter as a separate step

[0007] (ii) leaching the precious metal with a leach solution containinga thiosulfate-based lixiviant.

[0008] The present invention is based on the realisation that highlevels of precious metal recovery can be achieved on a cost-effectivebasis by carrying out precious metal oxidation and precious metalleaching as separate steps.

[0009] The material may be any material that contains precious metals.

[0010] The present invention relates particularly to materials in theform of ores and concentrates of the ores.

[0011] Preferably, the ores and concentrates are gold-bearing ores andconcentrates. The gold may be contained in oxidic or sulfidic ores.

[0012] In one embodiment treatment step (i) includes formingagglomerates of the precious metal-bearing material and an oxidant.

[0013] Preferably the agglomerates are formed by contacting the materialand a solution containing the oxidant.

[0014] More preferably this embodiment includes forming agglomerates ofthe material, a binder, and the oxidant.

[0015] More preferably the agglomerates are formed by mixing thematerial (such as an ore or concentrate of the ore) and the binder andthereafter contacting the mixture with a solution containing theoxidant.

[0016] Preferably, this embodiment includes curing the agglomerates.

[0017] Preferably the curing step is carried out in air for a period ofat least 24 hours.

[0018] The treatment step (i) may include forming agglomerates of theprecious metal-bearing material and the oxidant and a thiosulfate-basedlixiviant.

[0019] In another embodiment the treatment step (i) includes formingagglomerates of the precious metal-bearing material (with or without abinder) and thereafter contacting the agglomerates with a solutioncontaining the oxidant.

[0020] The treatment step (i) may include contacting the agglomerateswith a solution containing a thiosulfate-based lixiviant.

[0021] In a further embodiment the treatment step (i) includescontacting the material (without agglomerating the material first) witha solution containing the oxidant.

[0022] The treatment step (i) may include contacting the material with asolution containing thiosulfate-based lixiviant.

[0023] In each of the above embodiments, preferably the amount of thesolution of the oxidant is relatively small, typically between 10 and20%, more preferably, between 12 and 15%, by weight of the weight of theprecious metal-bearing material.

[0024] In each of the above embodiments, the treatment step (i) mayinclude treating the material with ammonia or an ammonium salt, such asammonium carbonate, to stabilise the oxidant.

[0025] The oxidant may be any soluble source of copper ions.

[0026] Preferably, the oxidant is selected from the group consisting ofcopper sulfate, copper salt, and ammonium complex of divalent copper.

[0027] The thiosulfate lixiviant may be any suitable soluble thiosulfatecompound.

[0028] Preferably the thiosulfate lixiviant is selected from the groupconsisting of sodium thiosulfate and ammonium thiosulfate.

[0029] The binder may be any suitable binder, such as a cement or anorganic binder.

[0030] The process of the present invention may be carried out under anysuitable pH conditions. In this connection, the applicant has found inexperimental work that the subject process can be operated over a widerpH range than prior art processes. Moreover, the applicant has foundthat the subject process is more flexible with operating pH than anumber of prior art processes and consequently pH adjustment may not benecessary—as is the case in these prior art processes.

[0031] The present invention may be carried out on a heap of preciousmetal-bearing material, such as gold-bearing ores and concentrates ofthe ore, by:

[0032] (i) passing the solution of the oxidant through the heap;

[0033] (ii) allowing the oxidant solution to drain from the heap;

[0034] (iii)passing the leach solution containing the thiosulfate-basedlixiviant through the heap; and

[0035] (iv) allowing the leach solution containing leached preciousmetal to drain from the heap.

[0036] The above sequence of process steps may be repeated as requiredto maximise recovery of precious metal from the heap.

[0037] The process may include a further step of processing the oxidantsolution that drains from the heap to recover the oxidant.

[0038] Preferably this step further includes recycling the oxidant tothe process.

[0039] The process may also include a further step of treating theprecious metal-bearing leach solution that drains from the heap torecover precious metal, such as gold, from the solution.

[0040] Preferably, this step also includes recycling thiosulfate-basedlixiviant to the process.

[0041] The present invention is not confined to process preciousmetal-bearing material in a heap and, by way of example, extends toother processing options such as continuously stirred tank reactors.

[0042] The process of the present invention can be applied to bothoxidic and sulfidic ores.

[0043] In the case of sulfidic ores, the conventional wisdom in theindustry is that such ores are refractory and that the sulfidic contentof the ores must be at least partially oxidised. However, it has beensurprisingly found by the applicant that the process of the presentinvention can be used to selectively oxidise the precious metal in theore while minimising or substantially avoiding oxidation of the sulphideore to sulfate.

[0044] The applicant has carried out experiment work on gold-bearingoxidic and sulphidic ores. This experimental work is discussed below.

[0045] The experimental work included the following basic process steps:

[0046] Step 1 Copper Pretreatment

[0047] A solution containing cupric ion (either as copper, copperdiammine or copper tetrammine) in a predetermined concentration wasprepared by dissolving a predetermined weight of anhydrous coppersulfate in a known amount of water. To this solution was added eitherammonia (so as to form copper tetrammine) or ammonium carbonate (AC) orbicarbonate (ABC) (so as to form copper diammine). This cupric solutionthus prepared was contacted with the ore for a fixed period beforeseparation by filtration (small scale) or natural draining (columns).

[0048] Step 2 Intermediate Wash (Optional)

[0049] If an intermediate wash was used, a predetermined volume of awash solution (either water or ammonia ˜0.87 M) was contacted with thefiltered/drained ore for a fixed period before furtherfiltration/draining.

[0050] Step 3 Thiosulfate Wash

[0051] The copper pretreated and (when performed) washed ore was thencontacted with a predetermined volume and concentration of eitherammonium or sodium thiosulfate solution for a fixed period beforefiltration or draining. Thiosulfate washing was repeated until little orno Au was detected in the collected filtrate. In some instances the orewas left in for extended periods between washes.

EXAMPLE 1

[0052] This example relates to small-scale leaching of high-grade oxideore (˜250 ppm Au)

[0053] The objective of this experimental work was to investigate atambient temperature the influence of:

[0054] (i) using CuSO₄ as a source of Cu²⁺ as opposed to differentammine systems (Cu—NH₃ to yield Cu(NH₃)₄ ²⁺ or Cu—AC to yield Cu(NH₃)₂²⁺ ),

[0055] (ii) using sodium thiosulfate rather than ammonium thiosulfate;and

[0056] (iii)exposure to air between sequential thiosulfate washes.

[0057] Table 1.1 summarises the series of experiments performed. TABLE1.1 Copper Pretreat Copper Intermediate Pretreat species washThiosulfate wash Compare copper species Cu-NH₃ Cu (NH₃)₄ ²⁺ Waterammonium thiosulfate Cu-AC CU (NH₃)₂ ²⁺ Water ammonium thiosulfate Cu-ACCu (NH₃)₂ ²⁺ Ammonia ammonium thiosulfate CuSO₄ Cu²⁺ Water ammoniumthiosulfate Compare thiosulfate type CuSO₄ Cu²⁺ Water sodium thiosulfate

[0058] The following is a summary of the experimental conditions.

[0059] (i) Wt of ore used (g, dry basis): 64

[0060] (ii) Copper pretreatment

[0061] wt. of copper sulfate (g):1.0(0.025M)

[0062] Total pretreat volume (m1): 250

[0063] Contact time with ore before filtration (min):15

[0064] No. of washes: 1

[0065] (iii) Intermediate Wash(when used)

[0066] Water:

[0067] Total Volume (ml): 300

[0068] Volume per wash (ml):100

[0069] No of washes: 3

[0070] Ammonia solution:

[0071] Total ammonia pretreat volume (ml): 250

[0072] Concentration (M): 0.87

[0073] No. of washes: 1

[0074] (iv) Thiosulfate wash

[0075] Volume per wash (ml):100

[0076] wt of ammonium thiosulfate(s)(g/100 ml wash)(when used): 3.7(0.1M)

[0077] wt of sodium thiosulfate pentahydrate(s)(g/100 ml wash)(whenused): 6.2 (0.1M)

[0078] Contact time of wash soln. with ore before filtration (min):5

[0079] No of washes : determined by Au content in filtrate (usually ˜8to 10)

[0080] Results are presented in FIGS. 1.1 and 1.2. These Figures areplots of cumulative % Au or Cu recovered in solution versus the numberof washes respectively. Where modifications to the usual sequence insequential leaching occurred these are highlighted in FIGS. 1.1 and 1.2.

[0081] Conclusion

[0082] In all cases with Cu pretreatment (of any form), the overall Auextraction level is either approaching or exceeding 90%. This suggeststhat high extraction levels may be achieved with the process of thepresent invention regardless of the form of the cupric ion.

[0083] The rate and extent to which copper desorbs mimics the trendsapparent in gold extraction.

EXAMPLE 2

[0084] This example relates to leaching of as received and agglomerated.low-grade oxide ore (˜6 ppm Au) using columns.

[0085] The most likely field application of the process of the presentinvention for low to moderate-grade ores would be as a heap or vatleach.

[0086] In order to investigate this process application, a series ofcolumns were fabricated using PVC tubing (D=50 mm, L=350−400 mm) andpacked with 1 kg of ore (dry weight basis) as illustrated in FIG. 2.1.Column leaching (which is a form of heap leaching) was then performedusing the process of the present invention and, to assess itsapplicability in the field, several trials of varying chemicalconfiguration were performed.

[0087] In general, columns were filled (to completely cover the bed) bypumping (from the bottom) or spraying (from the top) a predeterminedvolume of liquid (either pretreatment or leach). After soaking (usuallybetween ˜8 and 24 h), the liquid was allowed to drain and the ore rested(usually between 1-3 days) before the next soak and rest cycle wasbegun. Washings were collected and analysed for Au and Cu by AAS.

[0088] The column leach trials involved the use of two ore forms,generally referred to as:

[0089] (i) the as—received ore; and

[0090] (ii) agglomerated ore, where the ore was agglomerated with cementonly (usually using 5-6 kg of cement/t of ore.)

[0091] To determine the efficiency of column leaching using the processof the present invention (without the intermediate washing step) of alow grade oxide (˜6 ppm Au) ore by varying:

[0092] (i) the form of the ore:

[0093] agglomerated vs as-received (non-agglomerated);

[0094] (ii) the form of copper in pretreatment:

[0095] copper tetrammine vs copper sulfate; and

[0096] (iii)the amount of copper in the copper pretreatment step.

[0097] The following table (Table 2.1) summarises the experimentalmatrix performed. TABLE 2.1 Copper Pretreatment Form of Cu²⁺/ Weightconcentration Thiosul- (kg) equivalent of fate Leach Column (dry CuSO₄in Concentra- No. Ore type basis) g/l) tion (M) Compare the form ofcopper in pretreatment C1 Agglomerated 1 Tetrammine 0.1 (4 g/l) C2 asreceived 1 Tetrammine 0.1 (4 g/l) C3 Agglomerated 1 CuSO₄ 0.1 (4 g/l) C4as received 1 CuSO₄ 0.1 (4 g/l) Compare the amount of copper in thecopper pretreatment step C5 Agglomerated 1 CuSO₄ 0.1 (2 g/l) C6 asreceived 1 CuSO₄ 0.1 (2 g/1)

[0098] Results are presented FIGS. 2.1 a and 2.2 a. These Figures areplots of % Au recovered solution versus the cumulative weight ofrecovered solution for the two comparisons.

[0099] Conclusion

[0100] Comparison of the form of copper in pretreatment (Cu²⁺ vs Cu(NH₃)₄ ²⁺)

[0101] The best performed columns for Au extraction are those where theore was:

[0102] (i) pretreated with copper tetrammine (both agglomerated or asreceived ore; or

[0103] (ii) agglomerated and pretreated with CuSO₄ .

[0104] Comparison of the amount of copper in the copper pretreatmentstep:

[0105] Halving the copper concentration of the copper-sulfatepretreatment appeared to make little difference to Au extraction rate inthe as received ore but reduced extraction rate in the agglomerated oreby about half

EXAMPLE 3

[0106] This example relates to leaching of co-agglomerated low-gradeoxide ore (˜6 ppm Au) using columns.

[0107] In this example the ore was first pretreated with copper beforesubsequent thiosulfate treatment was performed. To reduce the number oftreatment steps and simplify operation in the field, it may be possibleto apply the required copper component by co-agglomerating it (inaddition to the cement) in the ore and thus avoid the pretreatment step.Field operation would then require only thiosulfate washing duringextraction. To this end a series of co-agglomerated ores were preparedwhere copper (as copper tetrammine) was added during agglomeration withcement.

[0108] Co-agglomeration was performed in the following manner:

[0109] Columns 7 & 8 Co-Agglomeration with copper.

[0110] To 3 kilograms of ore 18 g of cement was added. While this wasmixed 400 mls of a solution of 0.00156 moles/litre of copper as coppertetrammine was added.

[0111] Columns 9 & 14 Co-Agglomeration with copper and ammoniumthiosulfate.

[0112] To 3 kilograms of ore 18g of cement was added. While this wasmixed 200 mls of a solution of 0.00312 moles/litre of copper as coppertetrammine was added. In addition to this 200 mls of 0.26M ammoniumthiosulfate solution was added.

[0113] Comparing the extraction efficiency of ores co-agglomerated(besides cement) with either:

[0114] (i) small amounts of copper tetrammine (with and without an addedcopper pretreatment step); or

[0115] (ii) a combination of copper tetramnine and thiosulfate.

[0116] Leaches were performed in the manner previously described. Thefollowing Table (Table 3.1) presents the experimental matrix performed.TABLE 3.1 Copper Pretreatment Thiosulfate Weight Ore Form of Cu²⁺/ Leach(kg) Bed concentration Concentra- Column (dry L/D equivalent of tion No.Ore type basis) ratio CuSO₄ (g/l) (M) C7 Co- 1 6.4 None 0.1 agglomeratedwith copper tetrammine C8 Co- 1 6.6 CuSO₄ 0.1 agglomerated (1 g/l) withcopper tetrammine C9 Co- 1 4.9 None 0.1 agglomerated with coppertetrammine + thiosulfate C14 Co- 1 0.26 None 0.1 Agglomerated withcopper tetrammine + thiosulfate For comparison C1 Agglomerate 1 6.6Copper 0.1 tetrammine (4 g/l) C3 Agglomerate 1 7 CuSO₄ 0.1 (4 g/l) C11Agglomerate 1 0.26 CuSO₄ 0.1 (4 g/l)

[0117] Results are presented in FIG. 3.1. This Figure is a plot of % Aurecovered versus the cumulative weight of recovered solution.

[0118] Conclusion

[0119] The best-performed column (wide column) was that where the orewas co-agglomerated with copper tetrammine and thiosulfate.

[0120] Extraction behaviour decayed towards what appeared to be a limitof about 50%. To determine if the adsorbed copper level was a limitingfactor, the column was dosed with a treatment of copper ammine beforefurther thiosulfate washing was undertaken.. Although some subsequentincrease in Au extraction occurred, it appeared insubstantial andshort-lived. This suggested that, at this crush size, the ore might belimited to an extraction level of about 50-60%.

[0121] The treatments, where the ore was co-agglomerated with coppertetrammine alone (narrow columns C7, C8) showed no particular advantageand were abandoned after about 10 wash cycles. Co-agglomeration in widercolumns appeared to have the “initial kick” observed in small-scaleexperiments.

EXAMPLE 4

[0122] This example relates to leaching of co-agglomerated low-gradeoxide ore (˜6 ppm Au) using columns without using free ammonia.

[0123] The inclusion of ammonia or ammonium into the leaching system hasa beneficial effect during the early stages of the process of thepresent invention. However, in some environments the use of ammoniumthiosulfate may not be feasible because of its unavailability and theuse of free ammonia may also be restricted and sodium thiosulfate wouldbe used as a source of thiosulfate. However, if ammonium sulfate (asopposed to thiosulfate) is freely available it represents a source ofammonia/ammonium. On this basis, co-agglomerates were prepared wherecopper sulfate and ammonium sulfate were co-agglomerated to mimic thebehaviour of copper tetrammine.

[0124] Co-agglomeration was performed in the following manner:

[0125] Column 12

[0126] To 2.2 kg ore was added 11 gm cement (5 gm/kg). While mixing, 400ml of a solution containing 4 gm copper sulfate and 16 gm of ammoniumsulfate was added. (HIGH level)

[0127] Column 13

[0128] To 2.4 kg ore was added 12 gm cement (5 gm/kg). While mixing, 400ml of a solution containing 1 gm copper sulfate and 8 gm of ammoniumsulfate was added. (LOW level)

[0129] Table 4-1 presents the experimental matrix performed. TABLE 4.1Copper Thiosul- Pretreatment fate Weight Ore Form of Cu²⁺/ Leach (kg)Bed concentration Concen- Column (dry L/D equivalent of tration No. Oretype basis) ratio CuSO₄ (g/l) (N) C12 Co- 1 1.1 None 0.1 agglomeratedwith CuSO₄ and (NH₄) ₂SO₄ HIGH level C13 Co- 1 1.1 None 0.1 agglomeratedwith CuSO₄ and (NH₄) ₂SO₄ LOW level For comparison C14 Agglomerate 10.26 Tetrammine 0.1 (4 g/l) C11 Agglomerate 1 0.26 CuSO₄ 0.1 (4 g/l) C1Agglomerate 1 6.6 Tetrammine 0.1 (4 g/l) C3 Agglomerate 1 7 CuSO₄ 0.1 (4g/l)

[0130] Results are presented in FIG. 4.1. This Figure is a plot of % Aurecovered versus the cumulative weight of recovered solution.

[0131] Conclusion

[0132] With a co-agglomerated ore using high levels of Cu and ammoniumsulfate, Au extraction behaviour was similar to that of an oreco-agglomerated with copper tetrammine+thiosulfate

EXAMPLE 5

[0133] This example relates to leaching of co-agglomerated low-gradeoxide ore (˜6 ppm Au) in columns using a copper tetrammine made fromcopper sulfate, ammonium sulfate and sodium hydroxide and thiosulfate assodium thiosulfate.

[0134] Co-Agglomerated ores were made up as follows: Total CuSO₄ Ammon-Adjusted Ore ore (anhyd- ium with NaOH Na₂S₂O₃. Code wt Cement rous)sulfate to make 5H₂O : (kg) (kg/t) (kg/t) (kg/t) tetrammine (kg/t) 404 35 2 8 Yes 6.6 405 3 5 2 8 Yes 3.3

[0135]FIG. 5.1 presents % Au extracted (based on 6 ppm of Au in ore)versus weight or volume of recovered lixiviant per wash. Results for Aufrom the 404 and 405 are compared with previous best performing columnsthat had co-agglomerated ore with Cu-tetrammine+thiosulfateco-agglomerated ore with CuSO4+Ammonium sulfate (high)

[0136] Conclusion

[0137] The presence of copper tetrammine (made from either method) andthiosulfate in the co-agglomerated ore improves the initial rate ofextraction. Slight differences observed between C14 and X-404/X-405 maybe accounted for by differences in the thiosulfate concentration used inthe co-agglomeration step.

[0138] Based on the recovered solution analysis, the maximum extractionlevel was in the order of 50-60%.

[0139] At the end of the trials, residues from the best performingcolumns were fire assayed for Au and the extraction level calculated.This calculation indicated an extraction of 64-67%, a similar figure tothat determined on the as received ore from a cyanide-roll bottle test(˜56%). This suggests that the ore crush size may indeed be a limitingfactor.

[0140] To clarify this, a sample of as received ore was ring-milled andthen leached (in a high concentration thiosulfate, ammonia containinglixiviant system as per experiment 8). In this case, extraction levelrose to ˜77% confirming a limit on extraction due to crush size.

[0141] Many modifications may be made to the process of the presentinvention described above without departing from the spirit and scope ofthe present invention.

EXAMPLE 6

[0142] This example relates to leaching sulfide ores.

[0143] The copper pretreatment conditions were as follows:

[0144] copper tetrammine concentration (M): 0.025M

[0145] ammonia concentration (M): 0.235-0.435M

[0146] Total volume (ml): 250

[0147] The thiosulfate was conditions were as follows:

[0148] ammonium thiosulfate concentration (M): 0.1

[0149] volume per wash (ml): 100

[0150] Two ore/concentrates were examined: Kanowna Belle (X-136) andKCGM (X-133). The following effects were examined:

[0151] (i) premilling (by dry ring-milling for 5 minutes (RM))

[0152] (ii) varying the form of Cu²⁺ in the pretreatment step (Cu²⁺ CfCu (NH₃)₄ ²⁺)

[0153] Sequential leaches of pyrite concentrates were performed asdescribed above with the incorporation of various treatments. Thesetreatments included:

[0154] (i) leaving exposed to air or soaking in thiosulfate for extendperiods;

[0155] (ii) increasing the concentration of thiosulfate in the washsolution ; and

[0156] (iii) re-dosing ore with copper tetrammine.

[0157] Results based on solution analyses are presented in FIG. 6.

[0158] Conclusion

[0159] The highest Au extraction level was ˜50-60% using unmilledKanowna Belle (X-136).

[0160] Premilling appears to inhibit Au extraction although a greaterproportion of copper is adsorbed on the ore (60-70% cf 30-40%).

[0161] In all cases Cu adsorbed on the concentrate is readily desorbed.

1. A process for leaching precious metals from material containingprecious metals, which process includes the steps of: (i) treating thematerial by oxidising precious metal in the material into a form that isleachable in a subsequent leaching step; and thereafter as a separatestep (ii) leaching the precious metal with a leach solution containing athiosulfate-based lixiviant.
 2. The process defined in claim 1 whereinthe material is in the form of ores and concentrates of the ores.
 3. Theprocess defined in claim 2 wherein the ores and concentrates aregold-bearing ores and concentrates.
 4. The process defined in any one ofthe preceding claims wherein treatment step (i) includes formingagglomerates of the precious metal-bearing material and an oxidant. 5.The process defined in claim 4 wherein the agglomerates are formed bycontacting the material and a solution containing the oxidant.
 6. Theprocess defined in claim 5 includes forming agglomerates of thematerial, a binder, and the oxidant.
 7. The process defined in claim 6includes forming agglomerates by mixing the material (such as an ore orconcentrate of the ore) and the binder and thereafter contacting themixture with a solution containing the oxidant.
 8. The process definedin any one of claims 4 to 7 includes curing the agglomerates.
 9. Theprocess defined in claim 8 includes curing the agglomerates in air for aperiod of at least 24 hours.
 10. The process defined in any one ofclaims 4 to 9 wherein the treatment step (i) includes formingagglomerates of the precious metal-bearing material and an oxidant and athiosulfate-based lixiviant.
 11. The process defined in any one ofclaims 1 to 3 wherein treatment step (i) includes forming agglomeratesof the precious metal-bearing material (with or without a binder) andthereafter contacting the agglomerates with a solution containing theoxidant.
 12. The process defined in claim 11 wherein treatment step (i)includes contacting the agglomerates with a solution containing athiosulfate-based lixiviant.
 13. The process defined in any one ofclaims 1 to 3 wherein treatment step (i) includes contacting thematerial (without agglomerating the material first) with a solutioncontaining the oxidant.
 14. The process defined in claim 13 whereintreatment step (i) includes contacting the material with a solutioncontaining a thiosulfate-based lixiviant.
 15. The process defined in anyone of claims 5 to 14 wherein the amount of the solution of the oxidantis between 10 and 20% by weight of the weight of the preciousmetal-bearing material.
 16. The process defined in claim 15 wherein theamount of the solution of the oxidant is between 12 and 15% by weight ofthe weight of the precious-metal bearing material.
 17. The processdefined in any one of the preceding claims includes treating thematerial with ammonia or an ammonium salt, such as ammonium carbonate,to stabilise the oxidant.
 18. The process defined in any one of thepreceding claims wherein the oxidant is a soluble source of copper ions.19. The process defined in claim 18 wherein the oxidant is selected fromthe group consisting of copper sulfate, copper salt, and ammoniumcomplex of divalent copper.
 20. The process defined in any one of thepreceding claims wherein the thiosulfate lixiviant is selected from thegroup consisting of sodium thiosulfate and ammonium thiosulfate.